EP0441052A1

EP0441052A1 - Method for recovering zinc from zinc-containing waste materials

Info

Publication number: EP0441052A1
Application number: EP19900314001
Authority: EP
Inventors: Jan Arthur Aune; Inger Johanne Eikeland; Thor Pedersen
Original assignee: Elkem Technology AS
Current assignee: Elkem Technology AS
Priority date: 1989-12-22
Filing date: 1990-12-20
Publication date: 1991-08-14
Also published as: JPH0835020A; NO170032C; NO895250L; CA2032554A1; US5188658A; NO895250D0; NO170032B; PL288418A1

Abstract

A method for recovering zinc from zinc-containing materials, especially from zinc-containing material recovered from waste gases in metallurgical smelting processes. Particular zinc-containing materials are agglomerated together with a carbonaceous reduction material and optionally slag forming materials, and are supplied to a gas tight, closed, electrothermic smelting furnace (1) containing a molten bath kept at a temperature between 1200 and 1700°C in which the agglomerates are smelted and subjected to selective reduction and volatilisation of zinc and other volatile metals. An inert slag phase and optionally a liquid metal phase are tapped (8, 9) from the smelting furnace, and zinc and other volatile metals are recovered from the waste gas from the smelting furnace by condensation (6). In order to prevent reoxidation of metallic zinc, the particulate zinc-containing materials are agglomerated together with a carbonaceous binder which cracks at a temperature below 700°C and forms carbon black, and the temperature in the gas atmosphere in the smelting furnace is kept above 1000°C in order to maintain a volume ratio between CO₂ and CO in the gas atmosphere in the smelting furnaces below 0.3.

Description

The present invention relates to a method for recovering zinc from waste materials such as the dust recovered from waste gases in metallurgical processes such as steel production, but it can also be used for recovering zinc from ores, concentrates and other zinc-containing materials.
The dust which is recovered from waste gases in steel production will be referred to as EAF-dust, which is an abbreviation of Electric Arc Furnace Dust. The dust is precipitated from the waste gases in filter systems, as for instance bagfilters or other known filter systems. The particle size of the primary particles are usually within the range of 0.1 to 10 µm, but the primary particles are often partly agglomerated to somewhat greater particles. The EAF dust consists of complex oxides formed during the smelting- and refining processes as for instance oxides of Fe, Zn, Pb, Cd, Mn, Ni, Cu, Mo and other elements which are present in scrap iron.
The following Table 1 shows the chemical analysis of EAF-dust from four different steel producers. As can be seen iron oxides constitute the greater part, but the content of zinc oxide is also rather high. EAF-dust is a waste product which creates considerable problems and which up to recently has been impossible to dispose of or to use. Some of the oxides contained in the EAF dust are leachable in water and by deposition of EAF dust in landfills, these oxides may in the course of time be leached from the dust and can result in serious pollution of the environment.
A number of different methods for transferring EAF dust to such a form that it can be deposited without danger have been proposed. These methods however are not geared to recovering the valuable compounds in the dust. According to Norwegian Patent No. 160931 the dust is transferred to a form which makes it possible to deposite the product without any danger of pollution of the environment. In this process the dust is treated in a closed electrothermal smelting furnace in a multi-step process which also comprises the selective reduction and volatilisation of volatile metals. According to the Norwegian patent, zinc can be recovered by condensation. It has, however, been found that the yield of the zinc recovery in the condensation step is comparatively low due to the fact that a part of the volatilised zinc is reoxidised to zinc oxide in the gas atmosphere in the smelting furnace and in the zinc condensor.
The main reason for the reoxidation of zinc which occurs in the method according to Norwegian Patent No. 160931 is thought to be that due to partial reduction of iron oxides present in the supplied EAF dust, the CO₂ content in the gas atmosphere in the smelting furnace and in the channel between the smelting furnace and the condensor will be so high that the zinc vapour reacts with CO₂ to form ZnO and CO. The yield of zinc in the form of metallic zinc recovered in the zinc condensor according to the method of Norwegian Patent No. 160931 can therefore be as low as 55%.
It is an object of the present invention to provide a method whereby the yield of the recovered zinc can be substantially increased over the known state of the art.
Accordingly, the present invention provides a method for recovering zinc from zinc-containing materials (especially from zinc-containing material recovered from waste gases in metallurgical smelting processes) which comprises agglomerating particular zinc-containing materials together with a carbonaceous reduction material and optionally slag forming materials, supplying the agglomerates to a gas tight closed electrothermic smelting furnace containing a molten bath kept at a temperature between 1200 and 1700°C in which the agglomerates are smelted and subjected to selective reduction and volatilisation of zinc and other volatile metals; tapping of an inert slag phase and optionally a liquid metal phase from the smelting furnace; and recovering zinc and other volatile metals from the waste gas from the smelting furnace by condensation, characterised in that the particulate zinc-containing materials are agglomerated together with a carbonaceous binder which cracks at a temperature below 700°C and forms carbon black, and in that the temperature in the gas atmosphere in the smelting furnace is kept above 1000°C in order to maintain a volume ratio between CO₂ and CO in the gas atmosphere in the smelting furnaces below 0.3.
The temperature in the gas atmosphere is preferably maintained above 1000°C by heat radiation from the surface of the smelting bath. Preferably the supply of agglomerates to the smelting bath is regulated in such a way that 5 to 50% of the surface area of smelting bath is kept open and uncovered by agglomerates.
The temperature in the gas atmosphere can further be increased by blowing oxygen or oxygen enriched air into the gas atmosphere in the smelting furnace in an amount less than the stoichiometric amount necessary for combustion of the carbon black formed during cracking of the carbonaceous binder in the agglomerates.
According to a preferred embodiment the temperature in the gas atmosphere in the smelting furnace is maintained between 1100 and 1300°C and the volume ratio between CO₂ and CO in the gas atmosphere in the smelting furnace is maintained between 0.05 and 0.15.
Suitable examples of carbonaceous binders which can be used in the method of the present invention include resins containing a minimum of 50% by weight of carbon, such as tall oil pitch, petrol pitch etc. According to a preferred embodiment of the present invention tall oil pitch is used in an amount of 4-8% based on the weight of the agglomerates. Agglomerates are thereby achieved which have sufficient strength for ordinary transport in a smelting plant and which when heated in the smelting furnace, produce enough carbon black in order to keep the volume ratio between CO₂ and CO in the gas atmosphere in the smelting furnace at a sufficiently low level. Best results are obtained by using agglomerates containing 6 to 7.5% tall oil pitch.
By the method according to the present invention the agglomerates will, when they are charged to the smelting furnace, float on the smelting bath. The agglomerates are heated and while they are still in solid form, the carbonaceous binder in the agglomerates will be cracked and carbon black, H₂ and CH₄ will be formed. The volatile cracking products, together with part of the produced carbon black will escape from the agglomerates to the gas atmosphere above the smelting bath. In the gas atmosphere at least part of the carbon black will react with CO₂ which evolves during reduction of the metal oxides in the agglomerates, and form CO. The reaction between carbon black and CO₂ to form CO increases with increasing gas temperature, and at a gas temperature above 1000°C the volume ratio between CO₂ and CO can be kept so low that only a very small part of the zinc fume in the gas atmosphere will reoxidise to ZnO. In order to ensure a sufficient energy or temperature in the gas atmosphere and thereby achieve the best possible yield from the reaction between carbon black and CO₂, and for further increasing the temperature in the gas atmosphere, it is preferred to supply a controlled amount of oxygen or oxygen enriched air to the gas atmosphere in order to combust a part of the carbon black.
The part of the carbon black which is produced during cracking of the carbonaceous binder and which does not escape to the gas atmosphere, will immediately act to reduce oxides in the agglomerates and this reducing effect is adding to the reducing effect of the carbon containing reduction material present in the agglomerates. By using binders as described above, an extra effect is thus obtained which is not possible to obtain by using conventional binders such as bentonite etc. This also means that the amount of carbon-containing reducing materials in the agglomerates can be reduced.
As mentioned above, the agglomerates may contain slag forming materials. A preferred slag forming material is SiO₂-sand, and the amount is adjusted in such a way that a slag is produced having a sufficiently low viscosity that it can be tapped from the furnace. Wet dust is optionally dried before it is agglomerated. The agglomerates can favourably be preheated to a temperature below the cracking temperature for the binder, before they are supplied to the smelting furnace.
The invention may be carried into practice in various ways and one embodiment will now be further described by way of example with reference to the accompanying drawings, in which the single Figure shows an example of equipment for carrying out the invention.
In Figure 1, reference numeral 1 shows a gas tight electrothermic smelting furnace. The furnace can be of any conventional type, but it is preferred to use an electrothermic smelting furnace having a circular cross-section equipped with three carbon electrodes extending through the furnace roof in an absolutely gas tight manner. One such electrode 2 is shown.
The agglomerated raw materials containing a carbonaceous binder which cracks at a temperature below 700°C, are supplied to the furnace in a gas tight manner via a silo 3 and charging tubes 4 in such a way and in so many locations that the charge will float on the molten bath in the furnace and cover such a portion of the surface area of the molten bath that 5-50% of the surface of the molten bath remains uncovered by the charge. The charging tubes 4 can be arranged around the periphery of the furnace, about the electrodes, or centrally in the furnace.
The supplied agglomerates are heated in the furnace and melt at temperatures of 1200-1700°C, normally between 1300 and 1400°C. During reduction of the oxides a slag phase and a metal phase containing mainly iron, copper and nickel together with minor amounts of other metals present in the EAF dust are formed. The easily reducable and volatile elements, zinc and cadmium and a portion of the lead, are reduced and volatilised. If chlorides are present, they will also be volatilised, together with sulphur compounds and fluorides.
The gases which are formed in the smelting furnace are forwarded to a condensator 6 through a pipe 5. In the condensator 6, which can be of any conventional design, the volatilised zinc is condensed in a bath of molten zinc. If necessary a mechanical agitator 7 can be used to stir the bath. During contact with the molten zinc bath, the volatilised zinc will be condensed and will accumulate.
Molten zinc is continuously removed from the condensator in known manner. Minor amounts of zinc chloride, cadmium chloride and lead chloride which may floating on the top of the zinc bath and will be removed from time to time.
The remaining gas phase is optionally forwarded to an after-burner before any dust and other harmful components are removed from the gas in conventional ways.
As mentioned above a slag phase is formed in the smelting furnace. Below the slag phase there will exist a metallic phase containing iron, copper, nickel and minor amounts of other metals. The slag phase and the metallic phase are tapped from the furnace through two different tapping holes 8 and 9. Part of the lead oxide will be reduced to metallic lead in the smelting furnace and will accumulate in a separate metallic lead phase near the bottom of the furnace below the iron-containing phase. This lead is tapped at intervals through a tapping hole in the furnace bottom.
By the method according to the present invention it has been possible to obtain a yield, as metallic zinc, of more than 98% of the zinc which is volatilised from the agglomerates.
The invention will now be further illustrated with reference to the following non-limiting Examples.

EXAMPLE 1

A mixture of 87% EAF dust containing as main components 42.2% Fe₂O₃, 17.6% ZnO and 2.8% PbO; 8% silica sand (91% SiO₂); and 5% coke breeze (87% C) was briquetted under the addition of 7% tall oil pitch (79% C, 4% H (91% SiO₂); and 5% coke breeze (87% C) was briquetted under the addition of 7% tall oil pitch (79% C, 4% H and 15% O).
The briquettes were supplied to a lab scale smelting furnace having a freeze lining in its lower part. The furnace was equipped with two graphite electrodes. The temperature of the smelting bath in the furnace was adjusted to about 1400°C. The supplied briquettes which covered about 20% of the surface area of the smelting bath, were melted and formed a slag phase having a temperature between 1300 and 1380°C. The temperature in the furnace space above the molten bath was kept between 1110 and 1270°C by radiation heat from the uncovered part of the smelting bath. During the smelting of the briquettes, the iron oxides were mainly reduced to FeO which entered into the slag phase, while a minor amount was further reduced to metallic iron. Zinc- and lead oxide were reduced to the metallic state and volatilised.
Samples of the gas phase in the furnace were drawn by means of a water cooled lance and were continuously analysed to monitor the content of CO and CO₂. The volume ratio CO₂/CO was found to be within the range 0.1 to 0.2. A typical gas analysis showed 59.5 CO, 6.1% CO₂, 31.7% H₂, 1.2% CH₄ and 0.8% N₂. The off-gas from the furnace was forwarded to a zinc condenser containing liquid zinc operating at a temperature of 550°C. The volatilised zinc and lead in the gas phase were condensed in the condenser.
After the test was completed, the amount of metallic zinc formed in the condensator corresponded to 82% of zinc supplied to the furnace. In addition an amount of zinc oxide corresponding to 5% of supplied zinc was found. Of the volatilised zinc, 94.2% was thus recovered as metallic zinc. The produced slag contained an amount of zinc corresponding to 10% of the zinc supplied to the furnace. The slag contained 2.5% ZnO and 47.4% FeO.

EXAMPLE 2

A mixture of 87% EAF dust containing as main components 46.5% Fe₂O₃, 23.0% ZnO and 2.3% PbO; 8% silica sand (99% SiO₂); and 10% coke was briquetted under the addition of 7% tall oil pitch.
The briquettes were supplied to the same smelting furnace as described in Example 1 and the same temperature conditions were maintained in the furnace. The volume ratio CO₂/CO ws again found to be in the range 0.1 to 0.2. A typical gas analysis showed 78.9% CO, 7.9% CO₂, 11.2% H₂, 1.0% CH₄ and 0.7% N₂.
Zinc and lead were condensed in the zinc condenser. After the test run, metallic zinc corresponding to 83% of the zinc supplied to the smelting furnace and zinc oxide corresponding to 5% of the zinc supplied to the smelting furnace were recovered. Of the volatilised zinc, 94.3% was thus recovered as metallic zinc. In the slag, an amount of zinc corresponding to 8% of zinc supplied to the smelting furnace was detected. The slag contained 2.6% ZnO and 49.2% FeO.

COMPARISON EXAMPLE

A mixture of 87% EAF dust of the same composition as in Example 1, 8% silica sand (91% SiO₂) and 5% coke breeze (87% C) was briquetted with the addition of 7% bentonite (66% SiO₂, 28% Al₂O₃, 5% H₂O) as a binder and about 20% water.
A test smelting under the same conditions as in Examples 1 and 2 was run. The volume ratio CO₂/CO was now found to be in the range between 0.4 and 0.7 and a typical gas analysis showed 36.4% CO, 17.8% CO₂, 42.5% H₂, 1.7% CH₄ and 1.2% N₂.
Zinc and lead were condensed in the zinc condensor. Metallic zinc corresponding to 48% of the zinc supplied to the smelting furnace and zinc oxide corresponding to 41% of the zinc supplied to the smelting furnace were recovered. In the slag, an amount of zinc corresponding to 10% of the zinc supplied to the smelting furnace was detected. The slag contained 2.5% zinc and 47.1% FeO.
As it is evident from this comparison example, where a binder was used which did not contain carbon compounds which can be cracked, only 54% of the volatilised zinc was recovered as metallic zinc. Thus as much as 46% of the volatilised zinc was reoxidized to ZnO.
The examples thus show that by the method of the present invention, a very high reduction of the amount of zinc which reoxidizes to zinc oxide in the gas atmosphere can be obtained. The increased yield of recovered metallic zinc gives a much improved economy in the method for treating EAF dust, and further makes it possible to treat a number of other zinc-containing waste products and ores in an economically viable way.

Claims

A method for recovering zinc from zinc-containing materials which comprises: agglomerating particulate zinc-containing materials together with a carbonaceous reduction material and optionally slag forming materials; supplying the agglomerates to a gas tight, closed, electrothermic smelting furnace (1) containing a molten bath kept at a temperature between 1200 and 1700°C in which the agglomerates are smelted and subjected to selective reduction and volatilisation of zinc and other volatile materials; tapping an inert slag phase (8) and optionally a liquid metal phase (9) from the smelting furnace; and recovering zinc and other volatile metals from the waste gas from the smelting furnace (1) by condensation (6); characterised in that the particulate zinc-containing materials are agglomerated together with a carbonaceous binder which cracks at a temperature below 700°C and forms carbon black, and in that the temperature in the gas atmosphere in the smelting furnace (1) is kept above 1000°C in order to maintain a volume ratio between CO₂ and CO in the gas atmosphere in the smelting furnaces below 0.3.
A method as claimed in Claim 1, characterised in that the volume ratio between CO₂ and CO in the gas atmosphere in the smelting furnace (1) is maintained between 0.05 and 0.15.
A method as claimed in Claim 1 or Claim 2, characterised in that the temperature in the gas atmosphere is maintained above 1000°C by heat radiation from the surface of the smelting bath by regulating the supply of agglomerates to the smelting bath in such a way that 5 to 50% of the surface area of smelting bath is kept uncovered by agglomerates.
A method as claimed in any preceding Claim, characterised in that the temperature in the gas atmosphere in the smelting furnace is maintained between 1100 and 1300°C.
A method as claimed in any preceding Claim, characterised in that the temperature in the gas atmosphere is regulated by blowing oxygen or oxygen enriched air into the gas atmosphere in the smelting furnace in an amount less than the stoichiometric amount necessary for combustion of the carbon black formed during cracking of the carbonaceous binder in the agglomerates.
A method as claimed in any preceding Claim, characterised in that tall oil pitch is used as the carbonaceous binder in an amount of 4-8% based on the weight of agglomerates.
A method as claimed in Claim 6, characterised in that the tall oil pitch is used in an amount of 6 to 7.5% based on the weight of agglomerates.
A method as claimed in any preceding Claim, characterised in that the agglomerates are supplied to the furnace through charging tubes.
A method as claimed in Claim 8, characterised in that the charging tubes are arranged around the periphery of the furnace, arranged about the electrodes (2), or located in the centre of the furnace.
A method as claimed in any preceding Claim, characterised in that the agglomerates are preheated to a temperature below the cracking temperature of the carbonaceous binder before the agglomerates are supplied to the smelting furnace (1).